Method for treating floated solids

ABSTRACT

To prevent the flotation of a mineral containing an anionic collector reagent, the coated mineral is formed into a thick alkaline pulp containing another mineral having a greater affinity for the collector and an alkaline dispersant at a high concentration. The pulp is aged and then diluted and aerated. The mineral having the greater affinity for the anionic reagent floats; the other mineral is depressed.

United States Patent Mercade et al.

[451 *Jan. 18, 1972 METHOD FOR TREATING FLOATED SOLIDS lnventors:Venacio Mercade, Metuchen; Samuel R.

Weir, Long Branch, both of NJ.

Engelhard Minerals & Chemicals Corporation, Township of Woodbridge, NJ.

Assignee:

Notice: The portion of the term of this patent subsequent to July 8,1986, has been disclaimed.

Filed: Aug. 23, 1968 Appl. No.: 754,952

US. Cl ..209/3, 209/166 Int. Cl. ...B03b l/00, B03d 'l/06 Field ofSearch ..209/5, 166, 167, ll, 3

[56] References Cited UNITED STATES PATENTS 1,447,973 3/1923Feldenheimer ..209/5 3,331,505 7/1967 Mercade ....209/l1 3,454,1617/1969 Mercade ..209/167 Primary Examiner-Frank W. Lutter AssistantExaminer-Robert Halper Attorney-Melvin C. Flint ABSTRACT 10 Claims, NoDrawings METHOD FOR TREATING FLOATED SOLIDS This invention deals with animproved method for depressing particulate solid material containing ananionic collector reagent. The invention is especially directed to theflotation art and to the separation of minerals in slimed bulk flotationconcentrates obtained by negative-ion flotation concentration. Moreparticularly, the process of the invention represents an improvementover practices set forth in U.S.

Pat. No. 3,331,505 to Venancio Mercade.

BACKGRbUND OFTHE INVENTION The differential flotation of a mineral fromother mineral matter having similar flotation characteristicsisespecially difficult when the minerals are slimed. Fatty acidcollectors, for example, will float various oxide minerals from silicaand silicates, especially when well-dispersed pulps are used. However,the collector generally is not effective in selectively separatingslimed oxidized mineral from each other. Similarly, bulk'flotation ofslimed sulfide minerals may be made with fatty acids but selectivity isnormally not realized.

it has been suggested in U.S. Pat. No. 3,331,505 to Venancio Mercade toeffect the differential flotation of a slimed calcareous mineral,especially calcite, from a slimed noncalcare-' ous mineral in a fattyacid reagentized bulk float product by forming the bulk float into adilute pulp (.10 to 30 percent solids) containing an alkaline dispersantsuch as sodium silicate at a concentration of l to g./l. The dilute pulpis agitated until the froth is clear and nonslimy (about 10 minutes)while maintaining the pulp at a high temperature, e.g., 180 F. After thefroth or foam on the surface of the pulp becomes crystal clear, air isintroduced under the surface of the pulp. As a result, the calcareousmineral floats and is separated from the noncalcareous mineral whichremains dispersed in the pulp. Apparently reagent is transferred fromthe noncalcareous mineral to the calcite by the treatment with the hotdispersant solution. This may be attributed to the fact that thecollection of calcite is facilitated by the use of high temperature. Atany rate, under conditions of treatment employed in accordance with theteachings of the patent, heat is essential during the dispersanttreatment.

While unique separations have been effected by the process described inthe Mercade patent, the dispersant treatment of the bulk float to renderthe minerals separable by flotation without addition of anothercollector is undesirably costly for some beneficiation treatments. Thepulps are dilute, requiring a large volume of space and equipment. Thepower and heating requirements add further to the processing costs.

THE lNVENTlON An object of the invention is to provide an improvedprocess for treating a fatty acid reagentized bulk float product with adispersant (deflocculating agent) so that the solids in the bulk floatproduct may be separated by flotation without addition of a collectorreagent.

Another object is to provide an improved process for treating a solidcontaining a fatty acid collector reagent so that the solid is no longerfloatable.

Another object is to provide a novel method for separating slimedminerals having similar flotation characteristics.

Further objects and features will be apparent from the description whichfollows.

Briefly stated, in accordance with this invention, a semisolid or solidaqueous mixture of a particulate solid coated with an anionic reagentand a different particulate solid having a greater affinity for saidreagent is thinned by adding an a] kaline dispersing (deflocculating)agent without diluting the mixture to a solids content below 50 percentby weight. The

in accordance with an embodiment of the invention, the solid having thegreater avidity for the reagent may already be reagentized with theanionic collector when it is formed into the pulp, e.g., the solids inthe semisolid to solid mass may consist of a dewatered bulk floatproduct obtained with an anionic collector. Metallurgical results notheretofore considered possible have been realized when processinglowgrade slimed ore pulps by a primary bulk flotation with an anioniccollector, followed by dewaterlng of the bulk float, addition ofdispersant, aging, dilution-and aeration, all in accordance with thisinvention. By carrying out the bulk float before the secondarydifferential step, greater recoveries and concentration of valuedminerals may be realized than when attempts are made to recover theslimed mineral value initially by itself with a selective reagent.

Pursuant to another embodiment, the bulk float is a slimedultraflotation" froth concentrate obtained by the procedure describedfor example in U.S. Pat. No. 2,990,958 to Ernest W. Greene et al. Theprocess features the use of a reagentized finely divided solid additivesuch as calcite in a slimed fatty acid reagentized ore pulp. The addedoiled solid aids in the collection of the slimed selectively reagentizedconstituent of the ore and the two report in a bulk float concentrate.By applying the process to such a bulk float concentrate, the carrier isseparated from the ore constituent, usually as the float product in thedifferential flotation step. The carrier may be reused in subsequentultraflotation after being reoiled. The ore constituent thus separatedmay be'further processed.

The mineral having the greater affinity for the anionic reagent may beone that is added in unreagentized condition with a previously oiledsolid for the purpose of depressing the solid.

PRlOR ART v To the best of our knowledge, aging of high-solidsreagentized pulps in concentrated dispersant solutions has never beenemployed for any purpose, much less to permit subsequent differentialflotation without addition of selective reagents. We are aware that ithas been suggested in U.S. Pat. No. 2,811,254 to McGarry to agebulkphosphate-silica floats at high solids. However, the aging treatmentis not carried out in the presence of an added reagent such as ourdispersant. To the contrary, mineral acid, the reagent that removes theanionic collector in the McGarry process, is added after the aging stepand it is not present during the aging treatment. Moreover, the anionicreagent is removed on a nonselective basis in the McGarry process.Addition of a different type of collector (cationic) is then required toeffect the differential flotation in the secondary flotation treatment.ln contrast, our aging process, carried out in the presence of addeddispersant, permits differential flotation without addition of acationic collector since our dispersant treatment results in selectivedepression.

The process of the invention represents a substantial improvement anddeparture from practices set forth in said Mercade patent. A fraction ofthe equipment and space is required for the dispersant treatment andthe, agitating and heating equipment may be dispensed with. For example,when the dispersant treatment of the present invention is carried out ata preferred pulp solids level, e.g., 75 percent, the process requiresless than half the volume of equipment that would be used with theprocess of the patent at a maximum solids of 30 percent. With thehigh-solids pulps we employ during the aging treatment, the dispersantis present at a high concentration. When employed in quantity suflicientto thin the aqueous mixtures-as a result of deflocculation, the minimumconcentration of dispersant is at least 50 percent greater than themaximum concentration of dispersant employed in carrying out the Mercadeprocess. in most applications of our process, the dispersantconcentration is at least 500 percent greater than the maximumdispersant concentration in the Mercade process.

It was surprising and unexpected that selective flotation could becarried out after a pulp had been treated at the highdispersantconcentrations employed in our process. The prior art contains numerousreferences to the fact that minerals are depressed by alkalinedispersants when the dispersants are employed at concentrations that aresignificantly lower than those we employ. It would have been expectedthat all minerals including those previously collected would have beendepressed by the treatment and that differential flotation wouldtherefore not be possible. The beneficial results of the high-solidstreatment at high-dispersant concentrations were not predictable fromthose realized in carrying out the process of the Mercade patent. Theeffect of high temperature on the collector action of a fatty acid whichmight explain the operativeness of the Mercade process would not sufficeto explain why selectivity could be realized in carrying out our processwhich does not require the use of elevated temperatures. It issignificant that under the conditions of solids and reagent concentrations described in the Mercade patent differential separation wasnot realized at temperatures that may be employed in carrying out theprocess of this invention.

DETAILED DESCRIPTION The process is broadly applicable to the treatmentof mineral and/or nonmineral solids and is of special importance andvalue in the treatment of slimed, e.g., minus 200 mesh (Tyler) solidmatter.

The anionic collector may be any carboxy-containing reagent such as afatty acid or soap, especially a mixture of fatty and resin acids (orsoaps thereof). As is known in the art, anionic collectors usually alsoinclude hydrocarbon oils such as fuel oil and frequently also containemulsifying agents such as oil-soluble petroleum sulfonates. Examples ofanionic collectors include oleic acid, tall oil acids (refined orcrude), sulfo-oleic acid, soaps thereof, mixtures of the aforementionedwith each other or with hydrocarbon liquids such as fuel oil and/orpetroleum sulfonate salts. Especially good differential separations havebeen made from bulk floats containing a collector mixture of tall oilacids and an oil-soluble petroleum sulfonate.

The process of the invention is of especial importance in connectionwith differential flotation of minerals in slimed,flocculable-deflocculable bulk float concentrates containing anioniccollectors. In these processes, at least two mineral species arereagentized with an anionic collector or collector mixture. By applyingour process to such concentrates, the mineral species are separated andrecovered without the need to remove reagent from the bulk float as awhole and then add a different type of collector to effect thedifferential flotation. Separations not heretofore considered possiblehave been accomplished when the process of the invention was applied tolow-grade complex oxidized and sulfide ores. Especially good resultshave been realized separating slimed oxides of transition metals (e.g.,titanium dioxide such as anatase, tin oxide such as cassiterite, andmanganese dioxide such as pyrolusite) from calcite in finelydisseminated ores. After the bulk float was taken and the float productwas dewatered, dispersed, aged and aerated, the calcite floated and thetransition metal oxide was recovered in the flotation tailings. Forexample, manganese dioxide (pyrolusite) was recovered from a finelydisseminated ore containing calcite and siliceous gangue by producing abulk float with a fatty acid and then separating calcite by flotationfrom the pyrolusite by treating the bulk float in accordance with thisinvention. The separation and recovery of pyrolusite by this method isclaimed in a copending application, Ser. No. 754,951, filed of even dateherewith. Zinc sulfide and tin sulfide minerals have also been separatedfrom each other and gangue by similar treatment. The process is claimedin a copending application, Ser. No. 718,324, filed Apr. 3, l968 nowU.S. Pat. No. 3,454,161.

The present invention is also of value in the treatment ofUltraflotation froth products resulting from the flotation of slimedmetal oxide minerals, especially slimed anhydrous tetravalent metallicelements (such as anatase and cassiterite), with oiling(collector-coating) reagents including fatty acids and carrier particleswhich also have an affinity for fatty acids and are rendered floatablethereby. Minerals with surface characteristics similar to theaforementioned tetravalent metal oxides are rutile and zircon. Othermetal oxides include iron oxides and manganese oxides.

The process of the invention is applicable to the treatment ofUltraflotation froth products containing various types of auxiliarysolid or carrier flotation particles. Generally speaking, the auxiliarysolid flotation reagents are characterized by being very finely divided(minus 325 mesh) homogeneous particles that are different in compositionfrom the slimed pulp to be conditioned therewith. The carrier particlesare oiled by fatty acids and the oiled particles are floatable in theconditioned pulp. High-purity minerals, exemplified by minus 325 meshalkaline earth carbonate minerals, constitute one class of carrierreagent and are preferred in Ultraflotation processes for reasons ofeconomy. Alkaline earth carbonate particles include calcium carbonate,such as ground marble or calcite, magnesium carbonate, barium carbonateor mixed calciummagnesium carbonates. Nonwaxy organic solids, especiallynonwaxy polymeric solids, can also be used in Ultraflotation. Examplesof plastic carriers are polyethylene, polypropylene, polystyrene,polyvinyl chloride, polyamides and the like. Natural floaters, such astalc or the like, can also be used. Sulfur and fluorite are examples ofother carrier solids, Especially good results are obtained withmicron-size or slimed carrier particles, e.g., particles finer than 10microns.

The Ultraflotation bulk concentrates are obtained by the selectiveflotation of mineral particles from ore pulps fine enough to pass 200mesh (Tyler) screens. in some cases the pulps are fine enough to pass a325 mesh screen. The froths can also be obtained from the Ultraflotationconcentration of slimed pulps containing some coarse (e.g., plus 65mesh) particles.

ln carrying out the process of the invention, froths obtained by bulkflotation with negative-ion collector are thickened or dewatered,preferably by filtration, producing cakes of at least 50 percent solids.It is preferable to filter the cakes until the solids are within therange of 70 to percent since the use of high-solids cakes permits theformation of high-solids pulps which generally produce better results.There is no upper limit to the solids content of the cakes except forthe practical limitations involved in draining all of the water from thecakes. Since the water must be subsequently added to the filtered frothto dissolve the dispersing agent, it would be impractical to incur theexpense of removing all water. While the froths can be dried by the useof heat, the drying must be limited to the use of conditions which donot result in the destruction of the oils in the froths or in thecementation of the components. Decantation followed by filtration, orfiltration alone, are preferred methods for dewatering since they resultin the removal of pulp water which contains reagents that may interferewith subsequent operations. Also, when pulp water is removed bydecantation or filtration (alone or in combination with each other),entrained water-avid solids in the bulk float products may be removedsimultaneously.

Solids contents are determined by weighing the material before and afteroven drying at l75 F. for 1 hour and calculating as follows:

Percent solids Weight (original) weight (after drying) 7 W i a Thedefiocculating agent can be added to the dewatered pulp or filter cakeas dry reagent or as a concentrated solution. Incorporation of thedefiocculating agent to the solid or semisolid filter cake results in adistinct thinning or fluidization of the cake. Normally the fluidizedmass is very viscous and is frequently distinctly thixotropic. Sinceoiling reagents are also present, the thinned cake usually has a creamyappearance, similar to that of a milk shake. As mentioned above, thewater content of the deflocculated pulp must be restricted. Therefore,any water added to the filtered froth, separately or with thedefiocculating agent, must be limited. The addition to the filter cakeof the defiocculating agent as a concentrated solution results in a verysmall decrease in solids level.

To effect the deflocculation and resulting thinning of theflocculable-defiocculable solid or semisolid reagentized mass, thedispersant is added to the fiocculated semisolid mass in amount withinthe range of about to 50 lb./ton of solids (anhydrous dispersant basis).When this quantity of dispersant is added without reducing the pulpsolids level below about 50 percent, as required in our process, thedispersant concentration will be at least grams per liter of water inthe pulp; With most pulps, a dispersant concentration of at least 25grams per liter is required. Especially preferred is a dispersantconcentration of at least 50 grams per liter. By way of example, when 0sodium silicate solution (38 percent solids) is added without dilutionto a 73 percent solids flocculabledeflocculable aqueous pulp, the sodiumsilicate concentration (anhydrous basis) will be about 55 grams perliter of water in the pulp.

Stage addition of the defiocculating agent may be practiced, if desired.

After the defiocculating agent has been added to the dewatered froth,the resulting thick or viscous creamy pulp must be aged to permit thedefiocculating agent to act upon the constituents of the pulp in thedesired manner. During the aging, the pulp should contain at least 50percent solids, preferably 60 percent solids or more. Especiallypreferred is aging at a solids content within the range of 70 to 80percent by weight. The rate at which the defiocculating agent actsvaries directly with temperature. When the aging is carried out atambient temperature, times of at least 12 hours, preferably at least 18hours, are required. Aging can be carried out at temperatures up toabout 120 F. if desired, with corresponding reduction in time. There isno apparent advantage in our process to the use of very hot pulps.Prolonged aging is not detrimental.

The thick pulp may be maintained quiescent during aging or it may beagitated if sticking of the pulp to the sides of the vessel is aproblem. Opened or closed vessels may be used during the agingtreatment.

After the pulp has aged a suitable amount of time at the high-solidslevel, the pulp is diluted before the aeration and separationis carriedout. Dilution of a slimed pulp to a solids level of about 3 to 30percent, preferably about 5 to 10 percent, is recommended. Higher solidsmay be used when coarser pulps are employed. Dilution of the pulpresults in a decrease in the concentration of dispersingagent;additional dispersants or collectors are not added.

The diluted pulp is subjected to aeration in a suitable flotation cell.Aeration results in the flotation of oiled particles. Usually thecarrier particles float when a bulk Ultraflotation concentrate istreated. When reagent-avid solid such as calcite or fluorite has beenadded to depress a solid coated with an anionic collector, the formersolid is the one that floats. In the case of a bulk float consistingonly of ore minerals, the float product is a concentrate of the mineralthat has the greater affinity for the collector. For example, calcite isconcentrated in the float product when the bulk float has been obtainedby anionic flotation of an ore pulp containing calcium carbonate and anoxide of a transition metal such as anatase, pyrolusite, cassiterite.The mineral (or minerals) having the lesser affinity for the reagent orreagent mixture remains dispersed and depressed in the flotation cell.

The float products and tailings may be further processed by appropriatephysical and/or chemical means. Minus 325 mesh calcite is a preferredsolid when it is desired to depress a fatty acid reagentized oxide orsulfide mineral by the process of the invention. Other solids may beemployed. The dispersant and solid additive are added to a dewateredfloat product containing the mineral to be depressed, using sufficientwater and dispersant to provide a total solids content of at least 50percent and a suitable dispersant concentration, as described above.

The invention and some of its features and benefits will be betterunderstood from the following illustrative examples.

In the examples which follow all reagents are reported on an anhydrousweight basis unless otherwise indicated. Water that had been deionizedby treatment with cation-exchange resins and anion-exchange resins wasused throughout the mineral processing steps.

Examples l to VII illustrate the application of the process of theinvention to the separation and recovery of a carrier" mineral in aflotation process. The carrier used in these tests was minus 325 meshcalcite (Drikalite) obtained by grinding marble and classifying it to amean particle size below 5 microns. The Drikalite" is substantially purecalcium carbonate and has a GE. brightness value of 91 percent assupplied. This carbonate mineral was used as the carrier in the U1-traflotation concentration of colored (yellow-brown) anatase from kaolinclay with a fatty acid collector reagent in an alkaline pulp. As aresult of the presence of the anatase in the froth product of theUltraflotation beneficiation of the clay, the brightness of theDrikalite was reduced. Since the brightness of the treated frothgenerally varied inversely with anatase content, the brightness value ofthe treated froth was used in some instances to estimate purity of thecalcite float product.

EXAMPLE I The froth product that was used in the test was obtained froma commercial kaolin flotation operation, substantially as described inan article by Frank A. Seeton, Ultraflotation," Bulletin No. M4-B117,Denver Equipment Company. The clay charge was a Georgia kaolin from amine near McIntyre, Ga. The oiling agents employed included tall oil anda solution of neutral oil-soluble petroleum sulfonate in mineral oil(Calcium Petronate). Fuel oil has been employed during the flotation tocontrol the froth consistency. As supplied, the froth contained about 40percent solids and assayed 6.78 percent TiO The froth had a mustardlikeappearance and was composed of a mixture of the Drikalite carrier,colored anatase, adherent clay and oiling reagents.

The fresh froth from the flotation cell was washed free of clay by thefollowing steps. 0" sodium silicate solution was added to the froth inamount of 3 lb./ton solids and the mixture was diluted with deionizedwater to 1 percent solids. 0 sodium silicate solution contains 38percent solids and has a Na O: SiO mole ratio of 123.22.) The pulp wasmain tained quiescent to permit sedimentation. As a result, the clayformed a dilute clay-water suspension. Finer portion of the oiled frothsolids floated on the surface of the suspension. The coarser flocs inthe froth settled to the bottom of the claywater suspension. Theclay-water suspension was removed from the nonclay solids by siphoningthe suspension. The solids were again diluted to 1 percent and washed bysedimentation and siphoning. The procedure was repeated until the frothwas washed three times.

The froth solids remaining after the clay was removed were filtered,producing a solid filter cake containing 65 percent solids.

Q sodium silicate solution was added to the filter cake in amount of 1.3grams, corresponding to 26 lb./ton of cake solids, followed by 0.515grams dry sodium carbonate, corresponding to the use of 13 1b./ ton ofthis reagent. The dispersant concentration was calculated to be 55.7g./l. (anhydrous basis). The filter cake was deflocculated by theaddition of the sodium silicate and sodium carbonate, producing aviscous, thixotropic mass having the appearance of a thick milk shake.The deflocculated filter cake was agitated by a magnetic stirrer at lowspeed for 18 hours at ambient temperature. Since the dispersants wereadded in concentrated form, the solids content of the dispersed filtercake was only slightly less than 65 percent in spite of the fact thatthe dispersant had been employed as a solution.

After being stirred for 18 hours, the thick mobile mass was diluted topercent by adding deionized water and the diluted pulp was placed in al,000 g. capacity AirFlow flotation cell. The charge in the cell wasaerated and a float removed. The float was cleaned twice by flotationwithout addition or reagents.

The final float product was the calcite concentrate. This concentratewas dried, weighed and analyzed for TiO The TAPPl procedure was used tomeasure G.E. brightness.

The three machine discharges were combined to form the anataseconcentrate. This concentrate was flocced by adding alum and the floccedproduct was dewatered by decantation, dried and weighed. To determinethe TiO content, residual carbonate was destroyed by adding a 10 percentsolution of hydrochloric acid to the weighed material. The residue wasfiltered, washed and analyzed for TiO The results are summarized intable I.

TABLE I.-FLOTATION OF CALCITE FROM ANATASEIN CLAY FLOTATION FROTHCONCENTRATE l Calculated using fact that feed contained 90.1% CaCO Datain table I show that 98 percent of the calcite in the froth product ofthe kaolin flotation operation was recovered in the float product in theform of a concentrate containing only 0.34 percent TiO The data showthat 95.5 percent of the anatase in the froth product of the kaolinoperation reported in the flotation tailing, producing a concentrate of62.ll percent TiO grade.

These data therefore show that by aging the froth concentrate fromkaolin flotation with a concentrated dispersant solution at high solids,an excellent separation of the calcite from the anatase was effected byflotation without addition of oiling reagents or solvents. The resultsalso show that anatase was selectively depressed by the high solidsaging in the presence of the deflocculating agents.

EXAMPLE Il Example I was repeated with the principal exception that thefilter cake obtained after washing clay from the froth product of thecommercial kaolin flotation contained 73 per cent solids. (In Example Ithe cake was at 65 percent solids).) After dispersing the cake with 13lb./ton sodium carbonate and 26 lb./ton N" sodium silicate, thedispersant reagent concentration was 72.2 g./l., in contrast to exampleI in which the concentration was only 55.7 g./l.

Using the higher solids with more concentrated reagents and employing afroth assaying 8.34 percent TiO the calcite recovery was 98.1 percent.The calcite concentrate analyzed 0.07 percent TiO, and was thereforemuch purer than it was with the more dilute pulp. The brightness of thecalcite concentrate was 88.3 percent as compared to fresh calcite whichhad a brightness of 91 .0 percent. Using the higher solids pulp, thetailing assayed 77.52 percent TiO- representing 99.2 percent of the 'liOcontent ofthe froth from the kaolin flotation cells. In contrast. withthe more dilute pulp in example l, the

tailing analyzed 66.14 percent TiO at a similar weight recovery.

These results indicate that the anatase was more effectively depressedat the higher solids and that a sharper separation of calcite andanatase was effected when the pulp was aged with dispersant at 73percent solids than when the pulp was aged with the same dispersants at65 percent solids.

EXAMPLE lll Employing an Ultraflotation froth product substantially asdescribed in example I, and filtering the froth to 73 to 76 percentsolids, various deflocculating agents were added to the filter cakes andattempts were made to float the calcite from the anatase and clay by theprocedure employed in examples I and II. Employing 1.61 grams sodiumhydrosulfite or l.6l grams sodium sulfide per 200 grams filter cakesolids, aging for 18 hours, diluting to 5 percent solids and aerating,there was substantial selective flotation of calcite from the anataseand clay. Employing sodium carbonate as the sole dispersant in amount of13 lb./ton, there was some selective flotation of calcite but theresults were distinctly inferior to the results obtained with 25 lb./tonN sodium silicate alone or a mixture of 26 lb./ton N and i3 lb./tonsodium carbonate. Using 13 lb./ton tetrasodium pyrophosphate or 13lb./ton sodium hexametaphosphate, separation was effected although theresults were inferior to those obtained with the sodium silicate.

EXAMPLE IV A water-washed froth product from the commercial clayUltraflotation plant was filtered to 75 percent solids. To 200 grams ofthe filtered froth, 1.8 grams of lithium carbonate was added and themixture agitated. The mixture, which has the appearance of a thin milkshake, was covered and allowed to age at room temperature for 48 hours.Water was then added in amount sufficient to dilute the aged pulp to 5percent solids and the diluted pulp was agitated without aeration for 3minutes. The pulp was floated three times in a 1,000 gram Air- Floatflotation cell, producing a float product consisting of substantiallypure calcite and representing substantially all of the calcite in thefroth product from the Ultraflotation plant. The combined machinedischarge product or tailings contained clay and substantially all ofthe anatase in the froth product from the clay flotation plant.

EXAMPLE V Gray Georgia kaolin (a very fine particle size of sedimentaryhard clay having a distinct gray tinge) was subjected to Ultraflotationconcentration substantially as described in example I. An Ultraflotationfroth concentrate composed of an oiled mixture of Drikalite and anatasewas obtained. This froth concentrate was filtered. The filter cake (80percent solids) was deflocculated by adding dry sodium carbonate l3lb./ton), mixing, aging for 2 hours at ambient temperature, and thenadding N" sodium silicate (26 lb./ton) and aging for IS hours.

The aged pulp was diluted to about 5 percent solids, agitated in aFagergren cell without aeration for 5 minutes and then floated byintroducing air in the pulp. The flotation treatment was very effectivein bringing about a sharp separation since with only two cleanings an87.0 percent brightness calcite product was obtained as the floatproduct. The tailings was a Titania concentrate substantially free fromcalcite.

EXAMPLE Vl a. A sample of Georgia kaolin crude weighing 6,890 grams,corresponding to 6,000 dry clay, was blunged with 8,000 ml. deionizedwater, producing a 40 percent solids pulp. After being agitated for 10minutes, the pulp was dispersed by adding 240 ml. of a 5 percentsolution of 0" sodium silicate, corresponding to 4 lb./ton clay. Thepulp was then agitated for 10 minutes. To the pulp, 240 ml. of a 5percent solution of sodium carbonate was added, corresponding to 4lb./ton clay. The pulp was agitated for 10 minutes and degritted bypermitting the pulp to stand for 5 minutes and decanting the supernatantslip from the gritty sediment. The supernatant slip was thenfractionated on a centrifuge to produce a slip calculated to contain atleast 80 percent by weight of particles finer than 2 microns. The pH ofthe slip was 9.0. In a control flotation test, 2,280 grams of the slipof fractionated clay at 20.9 percent solids (500 gram dry clay) wasconditioned for flotation by adding the following: 100 grams Drikalite;30 ml. of a 5 percent aqueous solution of ammonium sulfate(corresponding to 6.0 lb./ton clay); 30 ml. of a 25 percent solution ofammonium hydroxide (corresponding to 3.0 lb./ton) and 90 drops of a50-50 mixture of refined tall oil (M-28") and mineral oil solution ofcalcium salt of petroleum sulfonate Neutral Calcium Petronate"). Thequantity of mixture corresponds to 9.0 lb./ton clay. Seventy-one dropsof petroleum hydrocarbon oil (Eureka M") corresponding to 8.0 1b./ton,was added. The pulp was conditioned for 20 minutes and had a pH of 8.9.

The conditioned pulp was floated in a Fagergren flotation cell, removinga froth for minutes. The froth was refloated three times and the machinedischarge products combined.

The brightness of the fractionated feed, flotation beneficiatedfractionated feed and bleached (zinc hydrosulfite) flotationbeneficiated clay were measured. A comparison of the results shows thatthe brightness of the clay was increased from 81.8 to 88.4 percent bythe flotation treatment and further brightened to 90.1 percent by thereducing bleach treatment. The machine discharge contained 452 grams dryclay, representing a 90.4 percent weight recovery.

b. Using the recovered, dried calcite from example II (88.3 percentbrightness) in place of the fresh Drikalite, the procedure of part (a)of this example was duplicated. After conditioning the pH of the pulpwas 9.0. The recovery and brightness of the flotation beneficiated claythat was conditioned with the reused carrier agent were substantiallythe same as the recovery and brightness using the fresh calcite carrier,thereby demonstrating that recovered calcite was a suitable substitutefor the fresh calcite.

EXAMPLE VI] A gray Georgia kaolin (80.9 percent G.E. brightness) wassubjected to Ultraflotation concentration, substantially as described inexample VI, substituting 75 grams powdered polyvinyl chloride for the*Drikalite carrier used in example VI. The plastic carrier was composedof particles within the range of I to 3 microns. Ninety percent of theclay was recovered. The beneficiated clay had an unbleached brightnessof 85.5 percent and a bleached brightness of 91.1 percent.

a. The forth product of the Ultraflotation concentration (calculated tocontain 84.1 grams solids) was allowed to stand, resulting in thesuspension of the solids in the froth concentrate. The liquid wassiphoned off and the froth solids washed by diluting them with water to5 percent solids and siphoning the water. This was repeated three timesin order to remove clay. The froth was then filtered to 75 percentsolids. To the filter cake, dry soda ash was added in amountcorresponding to 13 lb./ton of solids in the cake, followed by agitationand the addition of 26 lb./ton of 0" sodium silicate. The filter cake,which was fluidized by addition of the dispersant, was allowed to standfor 18 hours. The aged dispersion was mixed, aerated and floated threetimes without addition of reagents. The flotation tailings werecombined. Analyses of the flotation products showed that 85 percent ofthe plastic carrier was recovered in the float product. The flotationtailings included the colored impurities originally in the clay.

b. Similar results were obtained when the Ultraflotation froth productwas aged in the sodium silicate-sodium carbonate solution at 75 percentsolids for 2 hours at 180 F.

Thus, it has been shown that oiled carrier flotation reagents can beselectively floated froinan oiled slimed metal oxide addition, it hasbeen demonstrated that the reclaimed carrier can be recycled and be usedas the carrier in a subsequent Ultraflotation beneficiation operation.

Examples VIII and IX illustrate the application of the process of theinvention to the beneficiation of various slimed ores in which theminerals separated by the dispersant treatment of a bulk float mineralwere present in the ore per se. In Example VIII the bulk float containedan oxide of transition metal element (pyrolusite) and calcite. Asmentioned hereinabove, this specific embodiment of the invention isclaimed in our copending application Ser. No. 754,95 I. In Example IX,the bulk float was composed largely of sulfide minerals, principallyzinc sulfide and tin sulfide; these minerals were separated from eachother and recovered in accordance with principles of this invention.This specific embodiment is claimed in US. Pat. No. 3,454,161.

For purposes of comparison, results of conventional flotationbeneficiation processes are also given in examples VIII and IX.

EXAMPLE VIII A. Process of the Invention In accordance with thisinvention a concentrate assaying 52.4 percent Mn was obtained at anoverall recovery of 86 percent from a low-grade finely mineralizedmanganese ore from a deposit in the district of Corral Quemado, Chile.The ore assayed 23.1 percent Mn of which more than 90 percent waspresent as pyrolusite. Small amounts of manganese silicates were alsopresent. Gangue was predominantly calcite, quartz and silicates. The orealso contained barium sulfate and various carbonate minerals.

The manganese ore was crushed to minus 8 mesh and wetground in a pebblemill at 50 percent solids to 98 percent minus 200 mesh. To removesoluble salts, the ground ore was diluted to 10 percent solids withwater and the diluted pulp was allowed to settle. Supernatural liquidwas removed by decantation, leaving a washed pulp at about 25 percentsolids.

The pulp was dispersed by adding solid sodium carbonate in amount of 1.0lb./ton and then a hydrosol obtained by diluting 0 sodium silicatesolution to 5 percent adding a 1 percent solution of alum The hydrosolwas employed in amount equivalent to 8.0 lb./ton 0" sodium silicate and0.8 lb./ton alum. After addition of each reagent, the pulp wasthoroughly agitated.

After the pulp had been dispersed, ammonium sulfate was added in amountof 9 lb./ton. An alkaline collector emulsion was added, following whichthe pulp was conditioned for 5 minutes with a high-energy input in aDenver Sub A flotation cell. The emulsion was prepared in a WaringBlendor by mixing water with the equivalent of 2.0 lb./ton ammoniumhydroxide, 4.5 lb./ton of crude tall oil acids containing about 75percent fatty acid and 25 percent resin acids and 4.5 lb./ton CalciumPetronate." The emulsion contained about percent water. After theemulsion had been added, fuel oil (Eureka M) was added to thereagentized pulp in amount of 8.0 Ib./ton. The pulp was then conditionedfor 20 minutes.

A bulk float of manganese oxide and calcite gangue was obtained byaerating the conditioned pulp in a Denver Sub A flotation cell. Afterwithdrawing a froth for 7 minutes, the float product was cleaned twiceby reflotation without addition of reagents. In each cleaner flotation,the pulp was diluted to maintain adequate pulp level in the flotationcell. The flotation tailings from the bulk float operation werediscarded and the final float product (the concentrate of pyrolusite andcalcite) was dewatered by filtration, resulting in a filter cakecontaining about 70 percent solid. I

The filter cake was charged to a pug mill and deflocculated and thinnedby adding dry sodium carbonate in amount corresponding to 13 lb./ton,followed by addition of 26 lb./ton sodium silicate. The pug mill was inoperation while the dispersing agents were added. The resulting creamymass was held in a container at room temperature for 18 hours withoutagitation.

To separate the manganese oxide from the carbonate minerals in the mass,the dispersant-treated bulk float was diluted to about l0 percent solidsand dextrine was added to help depress the manganese. After conditioningthe pulp for minutes in a Denver Sub A flotation cell, air was admittedand a froth was withdrawn for 5 minutes. The froth was cleaned twice byflotation and the three machine discharge products were combined toproduce the manganese concentrate.

An X-ray diffraction pattern of the manganese concentrate indicated thatit was composed predominantly of pyrolusite and contained only traces ofcalcite. The concentrate represented 39.9 percent by weight of thestarting ore and analyzed 52.4 percent Mn, corresponding to a pyrolusiteproduct of about 85 percent purity. Overall recovery of manganese was86.0 percent.

B. Prior Art Treatment To illustrate the advantages of the process ofthe invention over prior art processes for concentrating manganese, themetallurgical results obtained in part A. of this example (process ofthe invention) were compared to results obtained when samples of orefrom the same deposit were concentrated by prior art flotation methods.

One of the prior procedures involved grinding the ore to 70 percentminus 200 mesh, followed by flotation of calcite from manganese using anoleic acid collector, fuel oil and quebracho. With an ore containing23.4 percent Mn, a concentrate analyzing 42.9 percent Mn (69.2 percentpurity) was obtained at a recovery of 69.8 percent. With a similarsample of the ore ground to 90 percent minus 200 mesh and flotation ofcalcite from the manganese with a fatty acid collector, dextrine todepress manganese and pine oil frother, the concentrate analyzed 43.5percent Mn and represented a 67.7 percent recovery.

A comparison of these results with the metallurgical results for theprocess of the invention which appear in part A. of this example, showsthat about 20 percent more manganese was recovered by the process of theinvention and that the manganese was recovered as a concentratecontaining about 9 percent more manganese. Thus, the process of theinvention was superior to the prior art processes with respect to boththe grade and recovery of manganese.

EXAMPLE [X This example illustrates separations of zinc and tin from acomplex zinc-tin Bolivian sulfide ore containing sphalerite, stannite(present as inclusions in the sphalerite), teallite (a solid solution ofthe composition PbS.SnS and tuffahlite (a zinc sulfide-tin sulfidemineral). Small amounts of cassiterite were also present. Gangueminerals include galena, pyrite, quartz and aluminosilicates.

A petrographic inspection of a representative sample of the oreindicated that the zinc and tin sulfide minerals were present in a stateof extremely fine dissemination.

Chemical assays of a representative sample showed the ore analyzedl4.37% by weight Zn, 2.03% Sn, 2.01% Pb, l4.54% Fe O 7.62% M 0 41.25%total SiO and 20.06% free SiO A. Process of the Invention A minus 8 meshsample of the Bolivian zinc-tin ore was ground at 50 percent solids in apebble mill to I00 percent minus 325 mesh. The ground ore pulp wasdiluted with water to l0 percent solids and the diluted pulp wasflocculated by adding sulfuric acid to a pH of 3.5. Supernatant liquidwas decanted and the thickened pulp was fluidized and dispersed byadding 10 lb./ton soda ash and 3 lb./ton O sodium silicate. Thedispersed pulp was conditioned for flotation by adding 6.0

lb./ton ammonium sulfate as a 5 percent aqueous solution and an emulsioncontaining water, l.0 lb./ton ammonium hydroxide, 6.2 lb./ton crude talloil and 6.2 lb./ton Calcium Petronate. The pulp was conditioned for 20minutes and floated in a subaeration-type flotation cell.

The bulk cleaner froth concentrate was cleaned three times byreflotation.

The tailings were discarded and the cleaned bulk concentrate wasfiltered. The filter cake which contained about 70 percent solids wasfluidized by incorporating l3 lb./ton soda ash and 26 lb./ton 0" sodiumsilicate solution. The fluidized cake was maintained in a closedcontainer for 20 hours.

The aged, dispersed filter cake was then diluted to about l5 percentsolids and aerated in the subaeration flotation machine. A froth product(the zinc concentrate) was withdrawn and recleaned twice by flotationwithout addition of reagents. The three tailings were combined, formingthe tin concentrate.

A summary of the overall results appears in table II.

TABLE ll PROCESS OF THE lNVENTlON-FLOTATION Data in table II show that68.5 percent of the tin in the ore was recovered in the final tinconcentrate containing 5.03 percent Sn. The recovery of zinc in the zincconcentrate was about 65.9 percent and this concentrate contained 45.68percent Zn by weight.

B. Prior Art Treatment A sample of the complex sulfide ore was crushedto minus 8 mesh (Tyler) and ground in a stainless steel rod mill at 50percent solids in the presence of soda ash (l5 lb./ton) to preventactivation of sphalerite and pyrite by iron from the mill. During the ginding, 0.6 lb./ton sodium cyanide and 2.0 lb./ton zinc sulfate wereadded to promote the deactivation of sphalerite. After grinding, the orewas diluted with water to 20 percent solids. The pulp was conditionedfor flotation of stannite by adding sodium hydroxide to a pH of 9.5 and0.025 lb./ton Z-l l" xanthate. A rougher tin-lead flotation was made.The rougher float was cleaned three times without addition of reagents,producing the final tin-lead concentrate and a combined cleaner tails.The rougher tailings were treated with 1.5 lb./ton copper sulfatepentahydrate to reactivate the sphalerite and 3.0 lb./ton lime for pHcontrol. The pulp was then conditioned for sphalerite flotation with0.075 lb./ton Z-l l" xanthate. A small amount of Dowfrother 250 wasadded. A second float was taken after addition of 0.075 lb./ton Z-l l"xanthate. The float products were combined to make the zinc rougherconcentrate. The flotation tailings were discarded. The zinc rougherconcentrate was treated with 2.4 lb./ton of lime and was cleaned twice,producing a zinc cleaner concentrate and a zinc cleaner tails product.

Metallurgical results for the test are summarized in table lll.

TABLE III Zinc Conc.

Tin Cone.

The data in table lll show that 75.1 percent of the Zn was recovered inthe form of a concentrate of 54.22 percent Zn grade by the conventionalsulfide process. However, only 5.5 percent of the Sn values wererecovered in the form of a concentrate of 3.65 percent Sn grade. Theresults therefore indicate that the zinc mineral responded well to theconventional sulfide flotation but that tin did not.

A comparison of the data in table ll (process of the invention) withdata in table lll (conventional sulfide flotation with depressant forsphalerite) shows that with the process of the invention the recovery oftin in the tailings was 68.5 percent, while in the conventional tinflotation process only 5.5 percent of the tin was recovered in the tinconcentrate. The data show also that the tin grade was higher when theseparation was made in accordance with the process of the invention.Zinc recovery and grade were slightly lower with the process of theinvention. Thus, the process of the invention resulted in a markedimprovement in tin recovery and grade without substantial sacrifice inzinc recovery or grade.

We claim:

1. In the beneficiation of an ore wherein an aqueous ore pulp containingfinely divided particles of alkaline earth carbonate mineral and finelydivided particles of a mineral which is an oxide of a transition metalis subjected to flotation in the presence of a fatty acid collectorreagent selective to said alkaline earth carbonate mineral and saidoxide of a transition metal, thereby forming a bulk float product whichis a fatty acid reagentized mixture of said minerals, an improved methodfor separating said reagentized minerals in said bulk float product fromeach other which comprises:

removing sufficient water from said bulk float product to form a masshaving a solids content of at least 50 percent by weight, incorporatingan alkaline deflocculating agent with said mass in amount sufficient tothin it while restricting the amount of water to an amount such that thesolids content of said mass is at least 50 percent by weight, aging saidmass at ambient temperature for at least l2 hours, diluting the agedmass with water to form a pulp, aerating the diluted pulp, therebyforming a froth product which is a concentrate of alkaline earthcarbonate mineral, and separating said froth from a tailing which is aconcentrate of said transition metal oxide mineral.

2. The method of claim 1 in which said alkaline earth carbonate iscalcium carbonate.

3. The method of claim 2 wherein said mass has a solids content withinthe range of 70 to percent during said aging.

4. The method of claim 2 wherein both of said minerals contain asubstantial portion of minus 325 mesh particles.

5. The method of claim 4 wherein said oxide mineral is pyrolusite.

6. The method of claim 4 wherein said alkaline deflocculating agentcomprises sodium silicate and wherein total concentration ofdeflocculating agent is at least 50 grams/liter.

7. The method of claim 6 wherein said carbonate mineral includes calciumcarbonate present as an ore constituent.

8. The method of claim 4 wherein said oxide mineral is Titania.

9. The method of claim 8 wherein said Titania is present in the form ofyellow colored anatase.

10. The method of claim 9 wherein said alkaline earth metal carbonatemineral is calcite which is added to an ore pulp containing said anatasefor the purpose of aiding the flotation of said anatase with said fattyacid collector.

Patent No. 3,635,337 Dated January 2 Inv'entofls) Venancio Mercade andSamuel R. Weir It is certified that error appears in theabove-identified patent and that said Letters Patent are herebycorrected as shown below:

Column 1 page 1 line [72] "Venacio Mercade" should read Venancio Mercade7 Column 5 line 26, "0" sodium silicate solution" should read "O" sodiumsilicate solution Column 6 line 46, "Petronate" should read Petronateline 71, "Q" sodium silicate solution" should read "0" sodium silicatesolution Column 7 line 62, "N" sodium silicate" should read "N" sodiumsilicate Column 8 line 70, "6,000 dry clay" should read 6,000 grams dryclay Column 12 Table III, add the following column headings Wt. Z Sn 2Zn Z Sn Zn Zinc conc. 20.23 0.85 54.22 9.2 I 75.1

Tin conc. 2.80 3.65 16.38 5.5 3.2

Signed and sealed this 27th day of June 1972.

(SEAL) Attest:

EDWARD M.FLETCHER,JR. ROBERT GOTTSCHALK Attssting Officer Commissionerof Patents FORM PO-105O (10-69) USCQMM-DC 50375-p5 w u.s. GOVERNMENTPRINTING ornc: was 0-365-334

2. The method of claim 1 in which said alkaline earth carbonate iscalcium carbonate.
 3. The method of claim 2 wherein said mass has asolids content within the range of 70 to 80 percent during said aging.4. The method of claim 2 wherein both of said minerals contain asubstantial portion of minus 325 mesh particles.
 5. The method of claim4 wherein said oxide mineral is pyrolusite.
 6. The method of claim 4wherein said alkaline deflocculating agent comprises sodium silicate andwherein total concentration of deflocculating agent is at least 50grams/liter.
 7. The method of claim 6 wherein said carbonate mineralincludes calcium carbonate present as an ore constituent.
 8. The methodof claim 4 wherein said oxide mineral is Titania.
 9. The method of claim8 wherein said Titania is present in the form of yellow colored anatase.10. The method of claim 9 wherein said alkaline earth metal carbonatemineral is calcite which is added to an ore pulp containing said anatasefor the purpose of aiding the flotation of said anatase with said fattyacid collector.